Classifications

• • C—CHEMISTRY; METALLURGY
  • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
  • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
  • C22B11/00—Obtaining noble metals
  • C22B11/08—Obtaining noble metals by cyaniding
• • C—CHEMISTRY; METALLURGY
  • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
  • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
  • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
  • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
  • C22B3/42—Treatment or purification of solutions, e.g. obtained by leaching by ion-exchange extraction
• • C—CHEMISTRY; METALLURGY
  • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
  • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
  • C22B11/00—Obtaining noble metals
  • C22B11/04—Obtaining noble metals by wet processes
• • C—CHEMISTRY; METALLURGY
  • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
  • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
  • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
  • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
  • C22B3/22—Treatment or purification of solutions, e.g. obtained by leaching by physical processes, e.g. by filtration, by magnetic means, or by thermal decomposition
  • C22B3/24—Treatment or purification of solutions, e.g. obtained by leaching by physical processes, e.g. by filtration, by magnetic means, or by thermal decomposition by adsorption on solid substances, e.g. by extraction with solid resins
• • C—CHEMISTRY; METALLURGY
  • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
  • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
  • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
  • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
  • C22B3/44—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes
• • Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
  • Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
  • Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
  • Y02P10/00—Technologies related to metal processing
  • Y02P10/20—Recycling

Definitions

• This invention relates to a method of recovery of platinum group metals (PGMs) from a solution or leachate containing the PGMs and base metals, and to the production of a PGM concentrate suitable as a feed stock to a PGM refinery.
• PGMs platinum group metals
• PGMs usually occur as discrete minerals, as dilute solid solutions in sulphide minerals, or are associated with silicates and/or chromitites.
• the processing of these materials consists of a combination of several unit operations, which ultimately produce individual PGMs - platinum, palladium, rhodium, ruthenium, iridium and osmium - plus gold and silver.
• PGM ores containing varying amounts of base metals such as copper, nickel and cobalt
• base metals such as copper, nickel and cobalt
• This flotation concentrate is dried and then smelted in an electric arc furnace where the PGMs are collected in the matte phase.
• the molten furnace matte is transferred to a converter where flux is added and air blown into the bath so as to remove the iron as a slag. During the blowing process, the majority of the sulphur in the furnace matte is converted to sulphur dioxide and emitted as a gas.
• the converter matte produced from the converting stage is either granulated or slow cooled.
• the granulated matte is milled and treated in a whole matte leach base metal removal circuit to separate nickel, copper, cobalt and iron from a PGM rich residue or concentrate.
• the base metals are sometimes refined to metal and in other plants produced as crude cathode and a semi pure nickel sulphate crystal, which contains iron and cobalt.
• the magnetic fraction is processed as described above, while the non-magnetic fraction, which is the larger mass fraction, is treated to recover nickel and copper as sulphates leaving iron in an insoluble fraction.
• the iron residue can contain up to 20% of the original PGMs emanating from the converter step.
• the high grade PGM concentrate is then treated in a precious metal refinery where the individual PGMs are separated and produced in pure metallic form.
• FIG. 1 A simplified flow sheet of the above process is shown in Figure 1.
• This conventional processing route for PGMs has its limitations. It is considered suitable for sulphide containing ores from which relatively low quantities of flotation concentrates are produced with high recoveries of PGMs that can be treated economically through a smelter. However, ores that have been oxidised produce greater amounts of flotation concentrates to achieve the same PGM recoveries and this can lead to uneconomic smelting operations.
• a process of recovering platinum group metals (PGMs) from a pregnant solution or leachate containing PGM values and base metals includes the steps of:
• the pregnant solution or leachate containing the PGM values and base metals is preferably a cyanide solution or leachate.
• the non-selective precipitation of the PGM values and base metals in step a) is preferably carried out by controlled reduction of the pH of the pregnant solution or leachate to within the range of about 0 to about 5, preferably about 1 to about 3, in particular about 2.
• the selective leaching in step b) is a pressure leaching step in which the PGM values remain in the residue.
• the PGM values are typically recovered from the residue by a fusion step or a further leach step to produce a concentrate rich in PGM values.
• the base metals are recovered from the leach solution, preferably through precipitation.
• the pressure leaching step may be carried out by first carrying out a caustic leach followed by an acid leach, in which case the PGM values may be recovered from the residue by, for example, a reduction leach step to produce a high grade PGM concentrate.
• the insoluble precipitate of step a) is fumed with acid, followed by selective leaching to form a leach solution containing both the PGM values as anions and the base metals as cations, and recovery of the PGM values on an ion exchange resin, which is preferably incinerated to form an ash rich in PGM values or eluted to recover the PGM values.
• the base metals are preferably recovered from the leach solution by precipitation.
• the insoluble precipitate of step a) is first calcined and then selectively leached to remove the base metals, producing a PGM rich residue or concentrate. Again the base metals are preferably recovered from the leach solution by precipitation.
• Any PGM values remaining in solution after the non-selective precipitation in step a) may be recovered by passing the solution through an ion exchange resin, as discussed above. Any base metals remaining in the solution may also be recovered, preferably by precipitation.
• Figure 1 is a flow sheet showing a prior art method of recovering PGMs
• Figure 2 is a flow sheet of an embodiment of the process of the present invention.
• Figure 3 is a flow sheet of an alternative embodiment of the process of the present invention.
• Figure 4 is a flow sheet of a further alternative embodiment of the process of the present invention.
• Figure 5 is a flow sheet of another alternative embodiment of the process of the present invention.
• Figure 6 is a graph that shows the effect of using NaHS over a range of pHs to precipitate PGMs and base metals
• Figure 7 is a graph that shows the effect of using H 2 O 2 over a range of pHs on recoveries of PGMs and base metals;
• Figure 8 is a graph that shows the effect of a combination of H 2 O 2 and
• Figure 9 is a graph that shows the effect of using sulphuric acid over a range of pHs on recoveries of PGMs and base metals from cyanide
• Figure 10 is a graph that shows the loading profile of platinum on the resins in a five-column pilot ion exchange circuit
• Figure 11 is a graph that shows the loading profile of gold on the resins in a five-column pilot ion exchange circuit
• Figure 12 is a graph that shows the effect of temperature of calcination on the leaching extraction of PGMs and base metals.
• Figure 13 is a flow sheet of a pilot campaign for treating a pH2 precipitate.
• the present invention provides a process to recover PGMs and base metals from a solution or leachate containing them, in particular a cyanide solution or leachate.
• the cyanide leachate is produced through a known cyanide leach process.
• a flotation concentrate or an acid pressure leach residue material bearing platinum group metals such as platinum, palladium, rhodium, ruthenium, iridium and osmium, other precious metals such as gold and silver, and various base metals such as copper, nickel and cobalt, is contacted with a cyanide leach solution to solubilise the PGMs and the base metals.
• the resultant cyanide leachate may, if desired, be concentrated to form a concentrate rich in PGM and base metal complexes.
• the cyanide leachate is acidified with diluted sulphuric acid in a pH range of 0 - 5,0 and conducted at a temperature in the range of ambient to 95°C.
• the acidification is preferably done using 1:1 sulphuric acid at around pH 2 and 45°C.
• the acidification step can take up to 6 hrs.
• the precipitation of the PGMs and base metals during acidification may be aided by controlled rates of addition of reagents such as sulphides in the form of sodium sulphide and sodium hydrogen sulphide or oxidants such as hydrogen peroxide and salts of peroxomonosulphuric acid (Caro's acid).
• reagents such as sulphides in the form of sodium sulphide and sodium hydrogen sulphide or oxidants such as hydrogen peroxide and salts of peroxomonosulphuric acid (Caro's acid).
• the precipitated metals are recovered from solution by solid-liquid separation. Any cyanide released as hydrogen cyanide can be recovered in a scrubber.
• the solution from the solid-liquid separation is passed through ion exchange columns where it is contacted with a resin that has functional groups that have stronger affinity for PGM complexes than the base metal complexes e.g. certain ammine groups.
• the PGM complexes bind to the resin in preference to certain base metal complexes of nickel, copper, cobalt and iron.
• the ion exchange columns are typically arranged in a "lead-trail" configuration. As the lead column becomes exhausted by the loading of the PGM and base metal complexes, the PGM complexes displace some of the base metals as additional feed is provided to the lead column. The trail column(s) captures these base metal complexes, as the case may be, and so a partition is achieved between the PGMs and the base metals. Eventually the base metals are displaced from the ion exchange resin and pass out with the raffinate.
• the resin may be removed and ashed to produce a PGM product.
• the resin may be eluted using an appropriate eluant to remove the PGM values.
• an ion exchange column loaded with a resin such as activated carbon is typically installed in the system.
• the residual base metal values are preferably precipitated from solution using reagents such as caustic soda and/or sodium carbonate and sodium sulphide.
• the precipitated base metal products can then be recovered from solution by solid-liquid separation.
• Alternative methods such as ion exchange, solvent extraction and electrowinning can also be used to recover the base metal values.
• the filtered precipitated residue from the initial step of acidification also referred to as the acidification residue, may be repulped using filtrate from a later filtration step.
• Sulphuric acid is typically added to the repulping step. Any off-gas is scrubbed to recover hydrogen cyanide (HCN), which is recycled to the cyanide leaching stage.
• HCN hydrogen cyanide
• the slurry from the repulping step is fed to a high-pressure vessel, typically a multi- compartment autoclave, where it is heated to a temperature of 140-220°C, preferably between 175 and 185°C. Oxygen is added counter currently to flush off any HCN into the vent from where it may be scrubbed.
• the autoclave is steam heated, directly or indirectly, and the pressure in the autoclave is maintained typically at 1500 kPa(g).
• the pulp density can be such as to yield a base metal concentrate varying from 10 to 180 g/I combined base metals, preferably 40 - 80 g/l.
• the autoclave retention time is between 1 and 6 hours, preferably 3 hours.
• the discharge from the autoclave is cooled to approx 85°C and flocculants such as Flocculant M351 added to aid settling before solid-liquid separation such as thickening and pressure filtration. After filtration, a solid product stream containing the PGMs and a filtrate stream are recovered.
• flocculants such as Flocculant M351
• the chloride conditioning process employs typically 5 to 10 g/l of chlorides at 90 to 95°C for approximately one hour.
• the raffinate from the ion exchange step can be treated by a variety of means to recover the copper. These include:
• the copper nickel ratio in the leachate subject to the copper nickel ratio in the leachate, metathetically exchanged for nickel-copper sulphide matte at temperatures between 70 and 180°C, preferably 130 - 140°C.
• the copper sulphide fraction can be treated to recover copper by a leach-electrowin process;
• the remaining nickel and cobalt in the raffinate stream after recovering the copper can be recovered by a variety of means. These include:
• the slurry from the high temperature atmospheric leach step is drained into a dilution vessel where it is diluted to typically 10Og/l free sulphuric acid.
• the leach is maintained at temperatures up to 95°C for typically 1 hr and then filtered using any suitable solid-liquid separation.
• the solid is washed before drying to produce a final concentrate rich in PGM values, typically assaying 15 - 85% PGM plus gold and silver, which may be refined in a conventional precious metal refinery to recover the PGM values.
• the filtrate and wash liquor after recovering the PGM concentrate may be recycled to the repulping stage of the acidification residue.
• pressure leaching of the acidification residue is conducted by first carrying out a caustic leach (sometimes referred to as 'oxydrolysis') followed by an acid leach as shown in Figure 3.
• This alkali-acid combination leach is conducted to destroy the cyanide complexes and solubilise the base metals for subsequent recovery, typically as described above.
• the leachate from solid-liquid separation is passed through a lead-trail ion exchange for the recovery of the precious metal values.
• the residue containing the PGM values may be reduction leached using a reductant such as formic acid to produce a high grade PGM concentrate suitable as a feed material to a conventional precious metal refinery.
• a further alternative embodiment, of the invention comprises a sulphuric acid fusion of the acidification residue followed by a chloride leach using reagents such as chlorine, hydrochloric acid and sodium chlorine or a combination of some or all the reagents ( Figure 4).
• the fusion/leach is conducted to produce a solution containing both the PGM values (as anions) and base metals (as cations).
• the PGM values are recovered by passing the solution through anionic ion exchange resin columns and followed by elution or alternatively ashing the resins.
• the base metal values are recovered by any suitable known method such as ion exchange, solvent extraction, precipitation or electrowinning.
• calcining of the acidification residue is conducted at a temperature between 250°C and 800°C, but preferably between 400-600°C, prior to an acid leach (Figure 5).
• the acid leach is conducted to solubilise the base metal values for subsequent recovery by any suitable known method.
• the PGM values remain in the residue and may be reductively leached to produce a high grade PGM concentrate.
• test work is presented below in a series of tests which have been conducted either using cyanide leachate, acidification solution or acidification residue.
• a test was conducted using NaHS to precipitate value metals from cyanide leachate at a raised temperature.
• a cyanide leach solution was prepared by calcining underground flotation concentrate at 400°C for 2 hrs in a muffle furnace and then leaching using 0,2% w/w sodium cyanide solution at 60°C for 48 hrs, maintaining the pH at 9,5 with hydrated lime.
• a cyanide leach solution was prepared by calcining a finely ground "open cut' flotation concentrate at 400°C for 2 hrs in a muffle furnace and then leaching using 0,2% NaCN solution at 60°C for 48 hrs, maintaining the pH at 9,5.
• a cyanide leach solution was prepared by calcining underground flotation concentrate at 400°C for 2 hrs in a muffle furnace and then leaching using 0,2% NaCN solution at 60°C for 48 hrs maintaining the pH at 9,5.
• Hydrogen peroxide H 2 0 2
• sulphuric acid was added to maintain the pH at set points between the ranges of 1-7.
• the reaction time was 120 min.
• Table 3 and Figures 7 and 8 show the effect of H 2 0 2 only and a combination of H 2 0 2 and Na 2 S 2 0 5 , respectively, over a range of pHs on recoveries of PGMs and base metals. It can be seen that with the exception of Au, good recoveries of PGMs and base metals were obtained at pH values at and below 3.
• a cyanide leachate of a PGM containing concentrate was acidified with 50% H 2 S0 4 over a range of pHs from 0 to 5.
• Cyanide leach solution was prepared by calcining underground flotation concentrate as per the previous examples.
• the alkaline cyanide leachate was warmed to 45°C and while gently agitating, 1:1 H 2 S0 4 was added slowly. Once the required pH was reached, the solution was allowed to stand while a precipitate formed. The pH was checked and readjusted to the target value.
• Flocculant M351 was added and allowed to stand for 3 hours without agitation.
• the concentration trends are shown in Table 4, and Table 5 shows the recovery of value material to the precipitate.
• PGM recovery to the precipitate was optimised at a pH of approximately 1.0. At pH2, the base metal residuals were at a maximum and then commenced redissolving. A similar trend was observed for the PGMs below a pH of 1.0.
• Figure 9 trends the precipitated component at various pH's.
• filtrates from Example 4 were subjected to ion exchange using a single quaternary ammonium ion exchange resin and a cation ion exchange resin.
• the first bed column had been kept in duty beyond its optimal cut point and was loading strongly held copper and cobalt complexes.
• Table 7 provides confirmation that a high grade PGM ash can be derived from the resin while Table 6 confirms that low raffinate PGM values can be derived employing the resin suite of IRA-402 and an activated carbon.
• the filtrate from the pH2 acidification step was passed through a series of ion exchange columns with the aim to saturate and analyse the resin in the lead column.
• Cyanide leach solution was prepared in a pilot plant run where calcined underground flotation concentrate, produced from a rotary kiln heated to 400-425°C and residence time of 2 hrs, was leached with NaCN in a 5 stage cascading circuit with a retention time of 100 hrs at a pH around 9,2 and temperature of 60°C. The solution was heated to 45°C before being pumped to a 3 stage cascading circuit with a retention time of 3-5 hours. Sulphuric acid solution of 500g/l H 2 S0 4 was added to the first tank to maintain the pH at 2.
• the filtered solution from the acidification step was then pumped at a flow rate of approximately 40 l/hr and 20°C into a series of ion exchange columns.
• the resin charges ofthe columns were as follows:
• the resin was equilibrated with 12 bed volumes of pH1 aqueous sulphuric acid prior to feeding pH2 filtrate.
• the load profiles for platinum and palladium are similar with the blend of IRA-402 and activated carbon employed as exchangers in the multi-column circuit.
• Figure 10 shows the load profile for platinum from pH2 leachates in the five-column lead-trail pilot ion exchange campaign.
• Figure 11 shows gold is not as strongly bound to the resin functional groups as platinum.
• Example 8 Recovery of PGMs from High Pressure / High Temperature Leach in a Two Stage Alkaline/Acid Medium followed by Ion Exchange.
• a series of tests were conducted by subjecting the pH2 acidification residue to a two-stage alkaline/acid leach at varying pressures and temperatures from 10-20 bar (1000-2000 kPa) and 150-200°C, respectively.
• Acidification residue was prepared as described in Example 4 and slurried with water in an autoclave and then steam heated to the required temperature. A weighed amount of caustic soda (NaOH) was added. The autoclave was pressurised to the set pressure and left to react for 2 - 4 hrs. A set volume of H 2 S0 4 was then added and again left for 1 - 2 hrs. The conditions for the two-stage leaches are shown in Table 12.
• the solutions generated by the pressure leaches were progressively passed through 10ml IRA 402 resin.
• the pH of the feed solutions varied between 0,49 and 5,83.
• a pH2 acidification residue was calcined at varying temperatures and then leached at high temperature and pressure.
• the calcine temperature varied from 250°C to 800°C and the calcined material was leached using dilute sulphuric acid at 6 bar (600 kPa) and 140°C for 4 hours.
• Figure 12 shows the effect of calcining temperature on the extraction of value metals. It can be seen that good separation is effected at a calcine temperature between 400-600°C where the PGM values remain in the residue and the base metal values essentially report to the solution. The results are summarised in Table 15.
• the pH2 acidification residue produced from the pilot plant testwork in Example 5 was weighed (800gm wet, 380-41 Ogm dry) into an autoclave and slurried with water. A known weight of concentrated sulphuric acid was added. The vessel was heated with steam to the target temperature and pressurised with oxygen. The reaction time was 4 hours. The residue from the pressure acid leach was then subjected to a concentrated sulphuric acid destruct (fusion) by blending with a measured volume of concentrated acid and heated with steam to 300°C for 1 hour. The leach slurry was diluted with water and filtered.
• This test run as a pilot campaign, illustrates the upgrade process for the precipitate produced from the pH2 acidification of cyanide solution.
• Example 6 The precipitate as produced in Example 6 was repulped in sulphuric acid, water and a lixiviant sourced from the second stage leach. The slurry was fed to the first stage leach.
• the flowsheet employed in this pilot campaign is shown in Figure 13.
• the first stage leach was an oxidative pressure leach with counter current oxygen flow.
• the vent gas was scrubbed to recover hydrogen cyanide that was not destroyed in the autoclave.
• the operating conditions for this stage leach were as follows:
• the mass reduction was only 80% (possibly due to only one compartment).
• the first stage residue was blended with sulphuric acid, leached at approximately 300°C in mild steel vessels and then quenched in water at 90°C.
• the conditions for the second stage of leaching were as follows:
• the PGM split between first stage leachate and the final concentrate is shown in Table 19.
• the final concentrate contained well in excess of 99% of the PGMs in the leach feed at a grade of greater than 80% total PGMs. Less than 0.1 % of the base metals in the leach feed reported to the final concentrate.
• the first stage leachate was conditioned with sodium chloride at approximately 15 g/l and then passed through pilot ion exchange columns filled with IRA-402 in the first, second and fourth columns and activated carbon in the third column.
• the raffinate values were typically 10 ppb for each element and the load distribution is given in Table 20.

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